The present invention discloses an innovative process for producing a titanium concentrate having contents of titanium and iron close to those usually found in ilmenite concentrates. Said material, herein called xe2x80x9cartificial ilmenitexe2x80x9d, is obtained by applying pyro and hydrometallurgical processes to the treatment of highly impure titanium ore, a class into which can be found in Brazilian ores, the predominant titanium source of which is anatase.
The processes for the production of titanium and its compounds are schematically shown in FIG. 1. Presently, there are only two ore sources commercially exploited around the world: rutile (TiO2) and ilmenite (FeTiO3). Several alternatives involving physical and/or chemical beneficiation processes for concentrating such ores have been proposed (HENRY, J. L. et all., xe2x80x9cBureau of Mines Development of Titanium Production Technologyxe2x80x9d, Bulletin 690, U.S. Bureau of Mines, 1987). The products obtained this way have essentially two destinations: direct production of metallic titanium and production of titanium pigment. The latter is presently carried out through the so-called sulphate and chloride processes, wherein the latter process is the main one and that with the highest growing potential.
Among the items shown in FIG. 1, the production of titanium slag (85% TiO2 equivalent) is of utmost importance due to the tact that this material can be directly used in the two presently available routes for the production of TiO2 pigment, by far the main application of titanium based raw materials. The following advantages can be mentioned for the production of slag: possibility of using titanium ore with a lower impurity content (ilmenite), utilization of the iron content of said ore through the production of pig iron, in addition to the already mentioned advantage of producing a material which can employed in the existing processes for the production of pigment. Estimates for the titanium slag consumption by the pigment industry show a present increase of up to 600,000 metric tons until year 2000 (ANON., xe2x80x9cTitanium Minerals 1994xe2x80x94A Complete Cost Analysisxe2x80x9d, AME Mineral Economics, 1994).
Brazil possesses approximately 20% of the known titanium ore reserves and the main concentration thereof can be found in alkaline pipes in the western region of Minas Gerais State and south of Goias State. The ore, that covers phosphate beds, is of complex nature, being the result of alterations in different types of rocks. The main difficulty in developing a process for recovering titanium amounts from this type of ore is the selection of single operations that show high efficiency and selectivity simultaneously. As a consequence, when trying to produce a concentrate rich in titanium the overall mass recovery falls dramatically, rendering the process uneconomical. Treatments of such kind have been tried in the past (PAI{overscore (A)}O, J. M. J. and MENDONA, P. A. F., xe2x80x9cProcess for Concentrating Titanium Oresxe2x80x9d, Brazilian patent PI 7507645; (PAX{overscore (A)}O, J. M. J. and MAGALH{overscore (A)}ES, G., xe2x80x9cProcess for Beneficiating Titanium Oresxe2x80x9d, Brazilian patent PI 7604532), resulting in alternatives of limited practical use, since mass recoveries have been comprised within the 5 to 10% range only.
The process here presented has been developed with the main purpose of overcoming the severe drawbacks found in the processes avaliable to date for the treatment of Brazilian anatase ores. As shown later; the solution for such a problem is obtained through by the utilization of iron amounts usually associated to titanium in the several anatase deposits found in Brazil.
The objective of the present process is the recovery on a titanium concentrate having a chemical composition similar to that of ilmenite; this concentrate being the intermediate raw material for the production of titanium slag. On the other hand, the process for producing slag is based on the carbothermic reduction of a part of the titanium contained in the orexe2x80x94leading to a material with a stoichiometry close to Ti3O5xe2x80x94and all the iron from the concentrate to FeOxe2x80x94contained in the slag phasexe2x80x94and metallic Fe. Other elements such as aluminium, silicon, alkaline and alkaline-earth metals are not reduced and thus constitute deleterious impurites that must be removed from the titanium raw material. Phosphorus is also an undesirable impurity due to the contamination caused in the resulting pig iron. As can be seen hereinbelow, the process developed makes it possible to obtain a high degree of removal of such impurities, thus providing concomitantly a large recovery of titanium and iron contained in the raw ore.
The process described in the present invention is based on the use of the following sequence of operations to the processing of titanium ores: disintegration, screening through 6 mm, crushing, classification, low intensity magnetic separation for removing coarse magnetite, attrition and slimes removal for discarding the minus 74 xcexcm (200 mesh) fraction, calcination in the presence of sodium carbonate (Na2CO3) and/or potassium carbonate (K2CO3), comminution of the calcination product to a particle size below 1 mm, leaching in alkaline medium (pH in excess of 10) and later acid leaching. Each of such steps is detailed below.
First of all, the raw ore is subjected to a disintegration step in a drum washer, followed by screening in a vibrating screen having a 6 mm (xc2xcxe2x80x3) opening, and crushing in a conical crusher. The screen operates in closed loop with the crusher, that is, only the  greater than 6 mm fraction is crushed. The screened material is classified in a spiral classifier which is adjusted for providing a product with a particle size above 48 mesh (300 xcexcm). The overflow of the classifier with a particle size of xe2x88x9248 mesh is discarded.
The classified ore is then subjected to a low intensity magnetic separation (approximately 800 Gauss), preferably in a wet drum separator. The magnetic fraction is discarded. The non magnetic fraction undergoes a slimes removal operation in an attrition cell with the purpose of removing clay-minerals particles that remain adhered to the surface of the anatase grains. The ore from the attrition step is classified in a spiral classifier adjusted in such a way that a 200 mesh (74 xcexcm) cut can be obtained. The overflow of the classifier, with a particle size ofxe2x80x94200 mesh, is discarded. For the purposes of the present invention, the material subjected to the above mentioned sequence of operations, that is, disintegration, screening, crushing, classification, magnetic separation, attrition and slimes removal, is called mechanical concentrate.
In the calcination step, the mechanical concentrate is mixed with sodium carbonate (Na2CO3) and/or potassium carbonate (K2CO3) and heated to a temperature within the 900 to 1300xc2x0 C. range, preferably 1100xc2x0 C., being kept at the desired temperature for a time period between 20 to 60 minutes, preferably 30 minutes, in an oxidizing environment. The amount of carbonate to be added, either as powder or in aqueous solution, depends on the content of Al2O3 and SiO2, ranging from 60 to 150 kg per metric ton of ore to be treated. The main purpose of this calcination is to produce soluble sodium compounds (Na2O.Al2O3, Na2O.P2O5, Na2O.SiO2 among others) at elevated temperatures, which are removed in the leaching steps that follow. No agglomeration or sintering was observed in the ore during calcination. Therefore, a simple disintegration of the calcined product is required in order to match its grain size with the next leaching step. In laboratory scale, this is carried out by compressing the calcined product in a rubber mantle, whereas in industrial scale it can be carried out through the use of conventional comminution equipment (hammer or jaw crusher) and care should be taken when regulating the operation of said equipment in order to avoid the generation of excessive fines. The calcined disintegrated product with a particle size below 1 mm feeds the leaching operations described hereinbelow.
The first leaching step is performed in alkaline medium (pH greater than 10) by using as a solvent any alkaline substance with a concentration lower than 50 g/L such as NaOH, KOH, Na2CO3 or K2CO3. In this step, the temperature is in the 60 to 90xc2x0 range, preferably 70xc2x0 C., and the time is 1 hour. After leaching, the material is filtered, and the liquor is reserved for the production of Al2O3, P2O5 or SiO5 and the recovery of the alkaline compound used. The solid residue of this step feeds the next step the acid leaching.
The product of the alkaline leaching is then subjected to leaching with HCl or H2SO4, preferably HCl, at concentrations ranging from 30 to 50 g of HCl per liter of solution, and preferably lower than 50 g/l. The temperature is in the 60 to 90xc2x0 C. range, preferably 70xc2x0 C., for a period of 20 to 240 minutes, preferably 60 minutes. After leaching, the pulp is filtered and the liquor is used in the recovery of rare earth compounds contained in the initial ore, and the solid, after drying, constitutes the final product, which is called artificial ilmenite. For the purposes of the present invention, the sequence of operations including calcinations, disintegration, alkaline leaching and acid leaching shall be referred to as thermo-chemical treatment.
The spirit and scope of the present invention may be fully understood based on the examples given below. It should be noticed that said examples are merely illustrative and shall not limit the process developed. The main steps of the process are shown in the schematic representation of FIG. 2.